TN295 



No. 9186 










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IC9186 



Bureau of Mines Information Circular/1988 




Rock Burst Research and the 
Coeur d'Alene District 

By Terry McMahon 




UNITED STATES DEPARTMENT OF THE INTERIOR 






CiUttJ Jtiito . . &</** if ¥*»*) 



Information Circular 9186 



Rock Burst Research and the 
Coeur d'Alene District 

By Terry McMahon 



UNITED STATES DEPARTMENT OF THE INTERIOR 
Donald Paul Hodel, Secretary 

BUREAU OF MINES 
T S Ary, Director 



Library of Congress Cataloging in Publication Data: 



^t^ 5 






McMahon, Terry. 












Rock burst research and 


the 


Coeur d'Alene 


District. 






(Information circular; 


9186) 










Bibliography: 42-45. 












Supt. of Docs, no.: I 


28.27:9186. 








1. Rock bursts— Coeur 


d'Alene Mountains 


(Idaho and 


Mont 


). I. Title. 


II. Series: Information circular 


(United States 


. Bureau of 


Mines); 9186. 


TN295.U4 [TN317] 


622s 


[622'.28] 




88-600025 



CONTENTS 

Page 

Abs tract 1 

Introduction 2 

Coeur d'Alene Mining District 2 

Initial discoveries and mining 2 

Geology of the district 3 

Rock, bursts 4 

Rock, bursts in other mining districts 6 

India 6 

Republic of South Africa 7 

Canada 7 

Other districts 7 

Rock burst mechanics 8 

Early investigations in the Republic of South Africa 8 

First phase 8 

Second phase 8 

Definitions of a rock burst 9 

Seismic energy sources 10 

Gravitational and tectonic stress 10 

Mining-induced stress 10 

Energy balance 10 

Energy release rate 13 

Conditions for a seismic event 13 

Unstable equilibrium 13 

Triggering stress increase 14 

Violent failure 14 

Release of seismic energy 14 

Failure mechanisms 16 

Intact rock 16 

Faulted rock 17 

Source locations and rock bursts 17 

At the face 17 

Ahead of the face 18 

On faults 19 

Coeur d'Alene rock burst research 19 

Microseismic method 19 

Early development 19 

Seismic source location 20 

Rock burst monitoring systems 22 

Rock failure mechanisms 22 

Dest ressing and preconditioning 22 

Pillar dest ressing 23 

Rock preconditioning 24 

First demonstration 24 

Second demonstration 28 

Stoping methods and mine geometry 31 

Overhand stoping 31 

Underhand stoping 33 

Underhand stoping demonstration 34 

Lucky Friday underhand longwall 38 

Conclusions 42 

References 42 

ILLUSTRATIONS 

1. Coeur d'Alene Mining District 3 

2. Structural evolution of Coeur d'Alene Mining District 5 



11 



ILLUSTRATIONS— Continued 



Page 



3. Split of total energy released with increase in number and decrease in size 

of excavation steps 12 

4. Increase in incidence and severity of rock bursts with increase in energy 

release rate 13 

5. Simple pendulum, illustrating unstable and stable equilibrium 14 

6. Mining-induced stress concentration ahead of face 15 

7. Stress-strain curve for perfectly-brittle material 15 

8. Stress-strain curves 16 

9. Simple spring-loaded sliding block model 17 

10. Types of shear fracturing in bedded quartzite ahead of stope 18 

11. Cumulative plots of noise source locations 21 

12. Holes drilled into pillar from stope below for destress blasting 23 

13. Seismic velocity contours as measured before and after destress blasting.... 24 

14. Vein preconditioning, first demonstration site, 7,700-ft level of Star Mine. 25 

15. Plan view of preconditioned zones, 7, 700-f t-level haulage lateral, and 

crosscuts to vein 26 

16. Preconditioning drill-hole patterns and loading, 7 and 10 crosscuts, 

7,700-ft level 27 

17. Vein preconditioning, second demonstration site, 7,900-ft level of Star Mine 28 

18. Preconditioning drill-hole patterns and loading, 8 and 12 crosscuts, 

7 , 900-f t level 29 

19. Pillar destress blastholes drilled from 9 crosscut, 7,500-ft level 30 

20. Preconditioning hole pattern, drilled from 7, 900-f t-level stope 31 

21. Generalized drawings of overhand and underhand stoping methods 32 

22. Model for simulated overhand and underhand stoping of section of vein 34 

23. Energy release rates for simulated stoping 35 

24. Underhand stoping demonstration site, 6, 700-f t level of Grouse vein, Star 

Mine 36 

25. Floormat construction details for cemented fill material 37 

26. Pillar destress blastholes drilled from stope below 6, 700-f t level 38 

27. Isometric view of Lucky Friday underhand longwall site 39 

28. Plan and cross section of Lucky Friday underhand longwall site 40 

29. Mining-induced stress -concentrations from simulated overhand (above 5,100-ft 

level) and underhand (below 5,100-ft level) stoping 41 





UNIT OE 


' MEASURE ABBREVIATIONS USED 


IN THIS REPORT 


ft 


foot 




in 


inch 


ft 2 


square foot 


min 


minute 


ft'lbf /ft 2 


foot 


pound (force) 


psi 


pound (force) per square 




per 


square foot 




inch 


fflbf/in 2 


foot 

per 


pound (force) 
square inch 


s 


second 
year 


ft/s 


foot 


per second 






h 


hour 









ROCK BURST RESEARCH AND THE COEUR D' ALENE DISTRICT 

By Terry McMahon 1 



ABSTRACT 

This Bureau of Mines report describes the rock burst problem and mea- 
sures being taken for its alleviation in the deep underground mines of 
the Coeur d'Alene Mining District in northern Idaho. The geologic fea- 
tures that contribute to rock bursting in that area are discussed and 
briefly compared with similar features in other burst-prone mining 
districts throughout the world. Early investigations to understand and 
control rock bursting are reviewed, and the mechanics of bursting are 
discussed in terms of the seismic events that produce them. Sources of 
strain energy, rock failure mechanisms, and the conditions necessary for 
the release of seismic energy are covered in detail. Rock burst 
research in the Coeur d'Alene District is described. Included are 
discussions of the microseismic method as both a research tool and as a 
rock burst monitoring system, rock destress blasting and preconditioning 
methods for rock burst control, and stoping methods currently being 
developed to reduce rock burst hazards. 



1 Mining engineer, Spokane Research Center, Bureau of Mines, Spokane, WA. 



INTRODUCTION 



Few hazards are feared by underground 
miners quite as much as the hazard of 
rock bursting. Rock bursts, unlike 
fires, explosions, and malfunctions of 
mechanical and electrical machinery, are 
beyond the control of the individual 
miner. Miners understand and are famil- 
iar with the safety hazards involved in 
the use of modern mining equipment and 
materials, and know that accidents can be 
prevented by training and safe work prac- 
tices. Unfortunately, this is not the 
case with rock bursting. The causes of 
rock bursts and effective measures for 
their control and prevention are, at 
best, only poorly understood. Further- 
more, rock bursts usually occur with 
little or no warning and thus leave 
individual miners feeling completely 
helpless and vulnerable against a phenom- 
enon capable of displacing hundreds or 
even thousands of tons of rock into their 
workplace. 

While mining activities have been prac- 
ticed for many millennia, rock bursting 
has only become a problem during the past 
100 yr. This is a direct consequence of 
the development of high-powered hoisting, 
pumping, and ventilation systems, and the 



use of powerful pneumatic drills and high 
explosives. These developments have made 
it possible to extend mine workings to 
greater depths than ever before, and thus 
into rock more highly stressed than ever 
before. 

The Coeur d'Alene Mining District is 
one area where severe ground control 
problems have developed. The incidence 
of rock bursting in district mines has 
grown over the past 40 or 50 yr as the 
mines have been extended to greater 
depths. Based on past experience, this 
problem will only become worse with con- 
tinued mining because the remaining ore 
reserves are in veins that go still 
deeper. Today, with the cost of mining 
higher than at any time in the past and 
because metal prices are often depressed, 
the added cost of rock bursting only 
serves to further threaten the economic 
survival of the district. 

This Bureau of Mines report has been 
prepared to provide an overview of the 
Coeur d'Alene District and the geologic 
environment that has resulted in its rock 
burst problem, and to review some of the 
specific research done by the Bureau to 
help alleviate the problem. 



COEUR D'ALENE MINING DISTRICT 



The Coeur d'Alene Mining District, 
located in the panhandle of northern 
Idaho, is one of many mining areas dis- 
covered in the Western United States 
during the latter half of the 19th centu- 
ry. Unlike so many other mining dis- 
tricts established during that same 
period, the Coeur d'Alene District's ore 
reserves were not exhausted after a few 
decades of intensive mining. Today, as 
the district enters its second century of 
mining, the deep vein ore reserves still 
extend to unknown depths. 

INITIAL DISCOVERIES AND MINING 

Discovery of the mineral deposits that 
started the rush to what would become 
Idaho's most famous and successful mining 
region was made in late 1883 by A. J. 
Prichard on the creek named after him. 



In January and February of 1884, the rush 
was on for placer gold, owing in large 
part to a brochure entitled "In the Gold 
Fields of the Coeur d'Alene," published 
by the Northern Pacific Railroad to stim- 
ulate business CL - J?.)•^ ^° e town of 
Murray, located in the northern part of 
the district, as shown in figure 1, was 
established within a few months, and by 
early summer placer mining was under way. 
Further prospecting led to the discovery 
of the lode deposits that were the source 
of the placer gold, and by 1885 several 
arrastras and stamp mills were in opera- 
tion to mill the gold-bearing quartz 
(3). 

^Underlined numbers in parentheses re- 
fer to items in the list of references at 
the end of this report. 




Fault movement 
Town 



FIGURE 1.-Coeurd'Alene Mining District. 



The vein deposits of lead and silver, 
for which the district would become world 
famous, were discovered by a disappointed 
gold prospector, Noah S. Kellogg, who 
wandered south and west of Murray in late 
1885 (4_). The discovery was made in Milo 
Gulch, a few miles south of the town that 
now bears Kellogg's name. One of the 
district's most famous mines, the Bunker 
Hill, was soon in operation, first as a 
glory hole and then as an underground 
mine. Within 5 yr, several other major 
silver-lead deposits had been located in 
the district and mining operations 
started. The towns of Wallace, Burke, 
Mullan, and what would become Kellogg 
were soon established, and by 1890, 



the district was being served by two 
railroads. 

GEOLOGY OF THE DISTRICT 

The host rock of the Coeur d'Alene's 
vein deposits is a regional sequence of 
Precambrian metasediraents known as the 
Belt Series. This sequence of rock 
extends from eastern Washington to west- 
ern Montana and underlies all of northern 
Idaho. The Belt Series consists of a 
thick conformable group of generally 
fine-grained quartzites and argillites 
(_5). The series is divided into six for- 
mations that range in thickness from 
about 1,000 ft to over 12,000 ft. In 



descending order from youngest to oldest, 
the formations are the Striped Peak, 
the Wallace, the St. Regis, the Revett, 
the Burke, and the Prichard. Mineral 
changes observed in all Belt Series rocks 
indicate that at least a slight regional 
metamorphisra has occurred during peri- 
ods of both high stress and high 
temperature. 

The geologic structure of the Coeur 
d'Alene District is extremely complex. 
Following deposition and lithif ication of 
Belt Series sediments, a system of prin- 
cipal folds is believed to have formed 
during a very early stage of deformation. 
This folding, probably the result of 
strong tectonic compression from the 
northeast and southwest, resulted in a 
series of synclines and anticlines. This 
stage of structural evolution is illus- 
trated in figure 2/4. As deformation con- 
tinued, deep-seated stresses may have 
become reoriented, subjecting the folds 
to a rotational stress that later became 
a shearing stress. These stresses caused 
a bowing and warping of the fold axes 
into a structural knot and also produced 
shear zones that would evolve into the 
major east-west fault zones of the dis- 
trict, as shown in figures IB and 2C. As 
shear zones evolved into continuous fault 
zones through the region, some of the 
strong, deep-seated tectonic stresses 
were probably released through the large 
fault displacements shown in figures 2D 
and 2E. The Osburn Fault is the most 
prominent fault of the district and has 
been mapped almost continuously for a 
distance of over 100 miles, from near the 
town of Coeur d'Alene, ID, to a point 
east of Superior, MT. The Osburn is a 
right-lateral, strike-slip fault with up 
to 16 miles lateral displacement and 
a nearly vertical dip through the 
district. 

Folding and faulting continued through 
the Larimide Orogeny and was accompanied 
by a period of monzonite stock intru- 
sions. This was also the period of prin- 
cipal ore deposition, which occurred in 
narrow, nearly vertical veins often 
extending to unknown depths (fig. 2F). 
The veins are generally located in linear 
bands along fractures and fault zones, 
and are generally found in the quartzites 



of the Belt Series as opposed to the 
argillites ^6). Continued stress and 
fault displacements have produced the 
tight folding and complex arrangement of 
fault offsets observed today both on the 
surface and in the deep mines of the 
district. 

The high tectonic stresses responsible 
for most of the district's complex struc- 
ture have left the area with an unusually 
high horizontal stress field. Measured 
horizontal stresses in district mines 
range from 1.5 to over 2.0 times the ver- 
tical stress (]_)• The vertical stresses 
encountered at the depths now being mined 
are often in excess of 6,000 psi. These 
factors — a hard, brittle rock mass in a 
very high stress field — produce a very 
high potential for violent rock failures. 

ROCK BURSTS 

Rock failures in underground mines may 
be referred to as either "rock falls" or 
"rock bursts." Rock falls, as the name 
implies, refer to falls of fractured, 
loose rock into mine openings. While 
potentially dangerous to miners and 
equipment and costly in terms of cleanup 
and lost production, rock falls are 
relatively nonviolent events. 

Rock bursts, on the other hand, involve 
the violent fracturing and explosive dis- 
placement of rock into mine openings. 
While a rock fall is generally a gravity- 
induced event, a rock burst results from 
the sudden and violent release of stored 
strain energy within the rock mass near a 
mine opening. When released, the strain 
energy becomes seismic, or shock, energy 
and travels through the rock as a seismic 
wave until the surface of an opening is 
encountered. The seismic energy, if of 
sufficient magnitude, is capable of frac- 
turing and displacing large volumes of 
rock. The magnitude of the seismic 
energy may vary over an extremely wide 
range, and as a result, the volume of 
rock displaced by a burst may range from 
no more than a few fragments to thousands 
of cubic yards. Small seismic events may 
be so weak that they are subaudible and 
thus require microseismic equipment for 
their detection; they do no damage to 
underground structures. Other seismic 



B 








Dobson Pass 
F 



Thompson Pass 
Fault 

sburn Fault 




Placer Creek Fault 



KEY 



fj Up war p 
Anticline 
Sy n c line 

Fault 



Monzonitic stock 
Major vein 
Tectonic force 



FIGURE 2.-Structural evolution of Coeur d'Alene Mining District, after Hobbs (5). 



events, termed "rock noise," may be aud- 
ible but still do no damage. At the 
opposite extreme are the large rock 
bursts, which have been recorded as hav- 
ing magnitudes as large as 2.7 on the 
Richter scale in the Coeur d'Alene Dis- 
trict and much higher in some other min- 
ing districts (8^). In the case of one 
Canadian mine, damage from a large rock 
burst hastened the complete closure of 
the mine (_9)» 

During the past 50 or so years, with 
mining depths in Coeur d'Alene mines 
reaching well below 2,000 ft, rock bursts 
have become common events. Until closure 
of the district's larger mines, bursts 
had become an almost daily event in one 
mine or another during periods of heavy 
mining activity (10). Most of these 
bursts would occur following mine blast- 
ing, when personnel were out of the 
stopes and away from potential burst 
areas. These bursts would often be rela- 
tively minor in magnitude and do little 
damage. In other cases, a rock burst 
might be preceded by rock noise in the 
form of a popping, cracking, or low 
rumbling sound, which would alert miners 

Ol> 

During the early years, miners often 
referred to rock bursts as "rock blasts" 
because of their explosive nature or as 
"air bursts" or "air explosions" because 
of the resulting airblast rushing through 



mine openings. Because of their violent 
and explosive nature, rock bursts fre- 
quently resulted in fatal or very serious 
injuries to anyone caught by the dis- 
placed rock. 

The extreme hazard that rock bursts 
present to Coeur d'Alene District miners 
has been underscored many times since 
June 4, 1941. It was then that two 
miners, William Mir and William Edison, 
became the first in a long list of rock 
burst fatalities in district mines (12). 
The two men were working on the 2900 
level of the Sunshine Mine when, at 
7:25 p.m., a major rock burst buried the 
men under tons of wall rock in the stope 
they had been working. The damage to the 
stope and amount of displaced rock was so 
great that rescue teams working continu- 
ously required 25 h to recover the first 
body and almost 3 days to recover the 
second. 

Since those first fatalities, there 
have been scores of other fatal accidents 
and serious injuries due to rock bursts 
in district mines. In addition, rock 
bursts have done hundreds of millions of 
dollars of damage to underground mine 
structures and equipment. As district 
mines are developed to greater depths, 
the potential for future rock burst prob- 
lems becomes greater because of the 
increased stresses encountered with 
increased depths. 



ROCK BURSTS IN OTHER MINING DISTRICTS 



The Coeur d'Alene District is not the 
only location where rock bursts have 
become a serious operational problem. 
The deep gold mines of India and espe- 
cially the Republic of South Africa have 
experienced severe rock bursts for many 
years, as have many other mining dis- 
tricts around the world. 

INDIA 

The first mining district to experience 
rock bursts was the Kolar goldfield in 
Mysore State of south-central India. The 
first reported burst occurred in the 
Oorganm Mine in 1898 at a depth of only 
960 ft, where gold was being mined from a 
narrow, steeply dipping vein (13). 



The ore vein is a belt of hornblende 
schist surrounded by granite and 
gneisses, which have been regionally met- 
amorphosed. The veins range in width 
from about 3 to 20 ft and have been 
interrupted by a series of faults, 
pegmatites, and dikes. These structural 
features have produced planes of weakness 
in the rock, which are believed to be the 
foci of the major rock bursts of the 
district. 

Rock bursts in the Kolar goldfields 
became a major operational problem as the 
mines went deeper. The deepest mine is 
now over 10,500 ft below the surface, and 
rock bursts with magnitudes of 4.5 to 5.0 
on the Richter scale have been recorded. 
Over the years, rock bursts have been 



responsible for numerous fatalities and 
injuries, heavy damage to mines (includ- 
ing the destruction of otherwise minable 
ore shoots), and even damage to surface 
structures located near the mine. In one 
instance, on October 22, 1937, a major 
burst was felt 18 miles away and damaged 
local seismometers to the point that 
local readings could not be obtained 
(14). 

REPUBLIC OF SOUTH AFRICA 

Gold mining on a major scale began in 
the Republic of South Africa in 1886 with 
the development of outcrops in the Wit- 
watersrand area of the Central Rand. The 
gold ore in this district occurs in a 
narrow conglomerate seam in quartzite and 
shale. Within 20 yr after the start of 
mining, when the mining depth was less 
than 1,000 ft, the first rock bursts or 
"earth tremors" began to be a problem. 

The gold-bearing conglomerate seam or 
"reef" is part of a sequence of geologi- 
cally stable Precambrian sediments, which 
dip from about 45° at the surface to 
about 25° at the 10,000-ft depths now 
being mined. The hanging wall consists 
of a thick sequence of quartzites, and 
the footwall is a sequence of shales. 
There are some offsets of the seam due to 
faults and intrusions by dikes, but the 
seam is essentially a planar structure. 
The actual gold-bearing portion of the 
conglomerate ranges from a few inches to 
several feet in thickness. 

During the early years in the Republic 
of South Africa, there was a dramatic 
increase in mining activity as well as 
associated seismic activity. Seismic 
records of the Union Observatory indicate 
only 7 tremors in 1908, but 233 tremors 
in 1918 when mining was at a depth of 
only about 2,000 ft. The rock burst 
fatality rate also rose dramatically dur- 
ing these early years, reaching a peak 
during the decade of 1926 to 1935. 
Records for the following 40 yr show a 
steady decline in overall fatalities but 
a nearly constant number of fatalities 
due to rock bursts (15). 



CANADA 

Deep, hard-rock mining in Canada has a 
history of rock bursting in districts 
such as Sudbury, Kirkland Lake, Timmins, 
and Red Lake, and also in the potash 
mines of Saskatchewan. Rock bursts were 
first reported at Sudbury in 1929. Since 
then, rock bursts at Sudbury, as well as 
in Canada's other districts, have been 
responsible for many fatal injuries and 
enormous damages. 

One example of a rock -burst-prone area 
is the Kirkland Lake District of Ontario 
where gold is mined from steeply dipping, 
narrow veins. The host rocks of these 
veins are Precambrian sediments, tuffs, 
and igneous intrusives that have been 
repeatedly folded, faulted, and intruded 
by dikes (16). This has produced a com- 
plex geologic structure similar to that 
of other burst-prone mining districts 
around the world. 

While several mines have operated in 
the Kirkland Lake District for many 
years, the Lake Shore Mine and the 
Wright-Hargreaves Mine have had more ser- 
ious rock burst problems. Rock bursts 
began at the Lake Shore Mine in the early 
1930's and shortly thereafter at the 
Wright-Hargreaves Mine. In 1964, a major 
burst at the Wright-Hargreaves Mine led 
to the closing of the entire mine (J^)« 

OTHER DISTRICTS 

Other mining areas where rock bursts 
occur include iron mines in Sweden (17) , 
the copper mines of Poland (18) , and var- 
ious underground mining districts in 
Europe. Some Australian and South Ameri- 
can mining districts have also reported 
problems with rock bursts. 

Underground mining is not the only 
excavation activity in which rock bursts 
have been a problem. In highway tunnel 
projects along western Norway's many 
fjords, rock bursts have occurred in the 
form of spalling rock caused by shear- 
fracturing of the tunnel surfaces. 
Investigations have revealed that these 
bursts are caused by both anisotropic 



stresses and large horizontal stress sides being tunneled (19). 
components within the steep mountain 

ROCK BURST MECHANICS 



Investigations into the source and 
cause of rock bursts and attempts to find 
preventative or control measures began in 
the Republic of South Africa shortly 
after bursting had become a serious 
problem. 

EARLY INVESTIGATIONS IN THE 
REPUBLIC OF SOUTH AFRICA 

The work in the Republic of South 
Africa is divided into two periods: the 
first beginning about 1908 and extending 
into the late 1940's, and the second 
beginning in the early 1950's and con- 
tinuing 'to the present. The first phase 
is known for its empirical approach to 
understanding the cause of rock bursts 
and trial-and-error methods of finding 
workable solutions. The second phase is 
known for its organized scientific 
approach to determining root causes and 
then designing preventative measures. 

First Phase 

The first phase of rock burst investi- 
gation began when the South African Gov- 
ernment appointed the Ophirton Earth 
Tremors Committee in 1908. Following its 
investigation, this committee concluded 
that the observed earth tremors were the 
result of the "shattering of support pil- 
lars." Their recommendation was to 
replace the solid reef pillars being left 
for roof support with packs of waste 
material. The Witwatersrand Earth 
Tremors Committee, appointed in 1915, 
concluded that the tremors were caused by 
"rock bursts due to sudden crushing of 
pillars," but also noted the "fracturing 
or settling of overlying strata" as an 
additional cause. This committee also 
recommended the use of waste packs or 
other artificial supports as a remedy. A 
third committee, the Witwatersrand Rock 
Burst Committee, was appointed in 1924. 
It classified rock bursts into two cate- 
gories: "strain bursts," which were 
small but numerous, and "crush bursts," 
which were less frequent but much larger 
(20). This committee also made several 



recommendations to help control the 
bursting problem. Among them were the 
use of artificial supports, the use of 
large shaft pillars for protecting mine 
shafts, and the use of longwall mining. 
The recommendation to use longwall mining 
methods indicates that the influence of 
mine geometry on abutment stress concen- 
trations was at least intuitively under- 
stood in the early years. Unfortunately, 
it was not until many years later that 
the longwall method was implemented and 
proved to be an effective rock burst 
control measure. 

The investigative work done during the 
early years was generally performed by 
the raining engineers working in the field 
and thus those who were closest to the 
rock burst problem. Understandably, 
their solutions were based on their 
observations and experiences and their 
own experimental methods, primarily trial 
and error. It is not too surprising then 
that often their recommendations to con- 
trol bursting were ineffective and 
occasionally even contradictory. As an 
example, some engineers advocated the use 
of solid backfilling for the artificial 
support of a mined-out area, while others 
recommended allowing the complete caving 
of the mined-out area. 

By the late 1940's, the South African 
gold mines were operating at depths 
greater than 8,000 ft and the rock burst 
problem was growing worse in both sever- 
ity and incidence. It then became appar- 
ent that available control measures were 
not effective, and if progress were going 
to be made, a more organized effort would 
be required. Furthermore, this effort 
would require a much more rigorous scien- 
tific approach. 

Second Phase 

In 1949, the second phase of rock burst 
investigations began with the formation 
of the Council for Scientific and 
Industrial Research (CSIR). Then, in 
1953, the South African Chamber of Mines 
assumed full responsibility for research 
on the rock burst problem. In organizing 



its plan of attack, the Chamber of Mines 
adopted four fundamental approaches to 
understand and alleviate rock burst phe- 
nomena. These were (1) observation and 
statistical analyses of rock burst occur- 
rences, (2) an intensive laboratory rock 
testing program, (3) an intensive inves- 
tigation of the seismicity of the mining 
areas, and (4) research into the applica- 
tion of mathematical-modeling methods for 
mine structures. 

The objective of the first approach, 
statistical analyses of rock burst occur- 
rences, was to establish relationships, 
if possible, between rock burst charac- 
teristics and various mining and geologic 
variables. Some of the important 
observations resulting from the work were 
that longwall mining methods reduce 
bursting, small or acutely shaped pillars 
produce higher incidences of bursts, 
burst activity increases significantly 
near the stope face, and bursting is 
often associated with certain geologic 
features such as dikes and faults (21 ). 

The program to test rock under labora- 
tory conditions was initiated to study 
the behavior of hard, brittle rock under 
high-stress loading conditions, to gain 
an insight into the deformation charac- 
teristics of such rocks, and to determine 
their failure mechanisms (22-23). Among 
the results of this testing was the 
discovery that hard, brittle rock, when 
slowly loaded in "stiff" testing 
machines, tends to exhibit strain- 
softening following the onset of failure 
(24-25) , rather than fracturing violently 
as brittle materials normally do when 
tested in "soft" testing machines. 

The investigation into the seismicity 
of the rock surrounding mines was the 
first to employ three-dimensional under- 
ground seismometer arrays. These permit- 
ted the accurate location of seismic 
events relative to the areas of active 
mining (26-27). By monitoring seismic 
activity underground, close to the 
source, it was found that the number of 
seismic events was far greater than the 
number of reported rock bursts. This 
led to the realization that not all seis- 
mic events result in rock bursts, 
and that the rock mass is much more 



seismically active around active mines 
than had been thought. 

Elastic-theory investigations were 
initiated to determine whether the 
response of the rock mass surrounding 
mines could be modeled numerically with 
models derived from the mathematical 
theory of elasticity. The displacements 
predicted by such modeling, when compared 
with measured displacements in the mines 
being modeled, indicated that most of the 
rock mass around mines did behave as an 
elastic material (28-30). Only the frac- 
tured and deformed rock in the immediate 
vicinity of mine openings, where inelas- 
tic behavior was exhibited, could not be 
modeled as an elastic continuum. 

These initial research efforts were 
only the beginnings of a continuing 
effort to obtain a basic understanding of 
rock burst sources and mechanisms and to 
find effective methods for their control 
and alleviation. Since these early 
efforts in the Republic of South Africa, 
rock burst research has continued not 
only there but in many other countries 
with mining districts where bursting has 
become a problem. The basic research 
conducted in more recent years has 
supplied a wealth of knowledge and expe- 
rience, and has focused on fundamental 
rock burst questions such as the source 
of energy, the mechanisms of rock failure 
that release this energy, and those min- 
ing variables and geologic conditions 
that make a rock burst more or less 
probable. 

DEFINITIONS OF A. ROCK BURST 

Many individuals and government agen- 
cies known for their experience and 
expertise in rock burst research have 
offered definitions of a rock burst. 
Some of these are 

A sudden rock failure characterized by 
the breaking up and expulsion of rock 
from its surroundings, accompanied by the 
violent release of energy (31 ). 

Damage to underground workings caused 
by the uncontrolled disruption of rock 
associated with a violent release of 



10 



energy additional to that derived from 
falling rock fragments (21 ). 

That phenomenon which occurs when a 
volume of rock is strained beyond the 
elastic limit, and the accompanying fail- 
ure is of such a nature that accumulated 
energy is released instantaneously (32). 

A sudden and violent failure of a large 
volume of overstressed rock, resulting in 
the instantaneous release of large 
amounts of accumulated energy (33). 

An instantaneous failure of rock caus- 
ing an expulsion of material at the 
surface of an opening or a seismic dis- 
turbance to a surface or underground mine 
(34). 

These definitions all point out the key 
characteristic of a rock burst: the sud- 
den release of energy in the form of 
violently expelled rock. The seismic 
energy released during a rock burst may 
range in magnitude from less than 0.5 to 
over 5.0, as measured on the Richter 
scale (35). The amount of failed rock or 
damage done by a burst will likewise vary 
over a wide range. 

In the Republic of South Africa's deep 
gold mines, a seismic event is considered 
a rock burst only when measurable damage 
is done. Based on past experience, this 
usually occurs when the seismic energy 
has a magnitude of 0.5 or more. Other 
fields, such as civil engineering, may 
have a definition that reflects higher 
sensitivity to rock bursting than does 
that of the mining industry. In such 
cases, any seismic activity induced by 
excavation may be considered a rock 
burst, even when there are no damages or 
injuries. 



Gravitational and Tectonic Stress 

In general, there are two natural 
sources of stress, gravitational and 
tectonic, that can deform a rock mass and 
produce stored strain energy. The gravi- 
tational stress at a given depth is 
simply a function of the rock density and 
depth below the surface. Tectonic 
stresses, on the other hand, originate 
from geologic stresses occurring on a 
regional scale and as such may contribute 
a larger stress component to the rock 
mass than the gravitational component. 
Tectonic stresses are often an important 
factor in burst-prone mining districts. 

Two burst-prone districts where tec- 
tonic stresses constitute a major compon- 
ent of the in situ stress field are the 
Coeur d'Alene District and India's Kolar 
goldfields. Horizontal stresses in the 
Coeur d'Alene District range from 1.5 to 
over 2.0 times the vertical stress (_7), 
and in the Kolar goldfields the ratio 
ranges from 1.6 to 4.0 (13). 

Mining-Induced Stress 

The stresses having the greatest influ- 
ence on the stability of rock surrounding 
a mine opening are those induced by 
excavation in the rock. The initial 
opening and enlargement of mine workings 
cause disruptions and consequent redis- 
tributions of the in situ stress field 
around the workings. Stress redistribu- 
tion takes place to compensate for the 
lost support of the mined rock and 
results in a concentration of stress 
about the opening. The total stress is 
then a combination of in situ and induced 
stresses. Together, these stresses 
strain the rock and produce stored strain 
energy. 



SEISMIC ENERGY SOURCES 



ENERGY BALANCE 



Since rock bursts are manifestations of 
seismic energy, it is important to con- 
sider (1) the sources of stress that 
strain a rock mass to produce stored 
strain energy and (2) the energy balance 
changes resulting from mining that may 
release the energy required to cause a 
rock burst. 



One of the results of early research 
was the realization that when an opening 
is made or enlarged, the balance of 
stored strain energy is upset (21 ). As a 
result of excavation, some energy is 
released when the newly formed opening 
surfaces converge slightly to adjust for 
the lost support of the removed rock. 



11 



Early research led to the conclusion 
that, for an unsupported opening, up to 
one-half of this released energy would 
become available as seismic energy. More 
recently it has been shown that this can 
only be true if the opening is made in 
one large step, i.e., removing all the 
rock at once. Where a mine opening is 
enlarged in many small steps (small rel- 
ative to the dimensions of the opening), 
the released energy can be accounted for 
almost entirely by the strain energy in 
the rock being removed by mining. Mining 
in small steps, as is usually the case, 
has thus been shown to be a stable activ- 
ity that does not result in the release 
of seismic energy (15). 

The energy change relationships oc- 
curring when a mine opening is enlarged 
from equilibrium state to another have 
been rigorously treated by Salamon ( 36 ) 
to determine the influence of the number 
of steps taken to make the enlargement. 
When the enlargement is made, energy 
becomes available from two separate 
sources. The first is the gravitational 
or potential energy W from the external 
and body forces doing work through dis- 
placement and deformation of the rock or 
volumetric convergence of the enlarged 
opening. The second is the stored strain 
energy U m contained in the volume of rock 
removed to enlarge the opening. The sum 
of these (W + U m ) is then the energy that 
becomes available as a result of mining 
and must be expended in some manner. 

Some of this energy will be expended by 
an increase in the concentrated strain 
energy U c stored in the rock mass ahead 
of the opening. If supports or backfill 
are used, then some of the available 
energy will be expended by support defor- 
mation W s . If it is assumed that the 
rock mass is an ideal elastic continuum, 
then none of the available energy will be 
dissipated by fracturing or inelastic 
deformation of the rock. With this 
simplification, the sum (U s + W s ) is the 
energy expended by enlargement of the 
opening. 

Since it is obvious that the energy 
expended cannot be greater than the 
energy available, and the strain energy 
U m removed with the mined rock is not 
available to do work on the rock around 



the opening or deform its supports, the 
inequality 



W > (U c + w s ) 
must hold, and since U m > 0, 
(W + U m ) > (U c + W s ). 



(1) 



(2) 



The difference in the energy available 
and the energy expended in inequalities 1 
and 2 is the excess or released energy 
W r . This difference must be expended in 
some manner, and from inequalities 1 and 
2, 



W r = (W + U m ) - (U c + W s ) > (3) 



and 



W r > U m > 0. 



(4) 



This shows that mining activity in an 
elastic continuum is always accompanied 
by the release of some energy, which must 
be expended in some manner. If the min- 
ing were done all at once, in one large 
step that removed all the rock instanta- 
neously, the sudden change would produce 
oscillations in the rock mass. These 
oscillations would then be dampened by 
minor imperfections in the rock, and 
kinetic energy W k would be dissipated 
as equilibrium was reestablished in the 
rock. 

Since no other means of energy dissipa- 
tion is available, inequality 4 becomes 



w r = u m + w k 



and from equation 3, 



Wi 



w 



(U c + W s ) > 0. 



(5) 



(6) 



Equation 5 indicates that the released 
energy W r is independent of the method of 
energy transfer, whether by the removal 
of rock U m or the dissipation of kinetic 
energy W k ; however, analytical computa- 
tions (36-37) have indicated that the 
relative values of U m and W k are heavily 
dependent on the number of steps taken to 
enlarge an opening. This is illustrated 
by the enlargement of both a circular 
tunnel and a spherical cavity from an 



12 



1 .0 




4 8 16 32 

NUMBER OF STEPS USED 
TO EXCAVATE TUNNEL AND CAVITY 



64 



FIGURE 3.— Split of total energy released with increase in number and decrease in size of excavation 
steps. 



initial to a final radius. The number of 
steps taken to make the enlargement 
ranges from 1 to 64. In both cases, the 
energy removed with the mined rock (U m ) 
increases, while the released kinetic 
energy W k decreases as the enlargements 
are made in progressivly more steps. 

These results are plotted in figure 3, 
and as can be seen, when a spherical 
cavity is enlarged in one step only, U m 
and W k each account for one-half of the 
released energy. As the number of steps 
is increased and the size of each step 
reduced, the energy released as kinetic 
energy W k is reduced and eventually 
approaches 0. The same trend holds true 
for enlargement of the circular tunnel, 
in that an increase in the number of 
excavation steps results in a reduction 
in the released kinetic energy. 

The results of this analysis, when 
applied to the energy changes due to min- 
ing, lead to the following equations: 



where the A symbol indicates that 
mining is done in very small steps. 



the 



Based on this work, Salamon reached the 
following conclusions: 

1. The work AW done by external and 
body forces through volumetric displace- 
ments is fully expended by the increase 
in strain energy AU C and deformation of 
supports AW S , when the change in mine 
geometry is very small. 

2. The energy AW r released by contin- 
ued mining is accounted for by the stored 
strain energy AU m in the rock removed by 
raining. 

3. There is virtually no kinetic 
energy AW k released by mining activity, 
and thus mining in small steps is a 
quasi-static, stable process and cannot 
be the source of seismic energy. 



AW = AU C + AW S , AW r 



AU r 



AW k = 0, (7) 



13 



As stated previously, the analysis and 
conclusions given above pertain to an 
ideal elastic continuum. As such, the 
results provide a limiting case when 
applied to mining. In an actual mining 
environment, the rock immediately sur- 
rounding an opening is usually highly 
fractured and thus exhibits inelastic 
behavior. The effect of inelastic behav- 
ior is to increase the volumetric dis- 
placement, which in turn increases the 
available energy W and the released 
energy W r . While not all of the released 
energy can then be accounted for by the 
energy removed with the mined rock (U m ), 
most of it can be, according to Salamon. 
Furthermore, the amount of energy that 
may become seismic energy is much less 
than previously thought. 

ENERGY RELEASE RATE 

"Energy release rate" is a term given 
to the calculated rate at which energy 
becomes available per unit wall-rock area 
mined, as a result of mining-induced dis- 
placements and stress changes. The basis 
for the calculation method is the fact 
that the spatial rate at which energy is 
released during mining is heavily depen- 
dent on mine geometry and the mining 
sequence taken to reach the final geome- 
try. When a particular stoping sequence 
induces high stresses and large displace- 
ments, the rate of energy release will be 
high. If the energy released is greater 
than can be dissipated by deformation and 
nonviolent fracturing in the rock, the 
excess energy will be released violently 
and may produce a rock burst. The calcu- 
lation method does not assume a threshold 
value of energy released above which rock 
bursting will occur but has been empiri- 
cally related to increases in both the 
incidence and severity of rock bursting. 
The energy release rate can be controlled 
to some degree by the systematic sequenc- 
ing of mining steps and, as a mine design 
and analysis tool, has been used exten- 
sively for mining tabular ore bodies in 
the Republic of South Africa where it was 
developed (21 ). 

The relationship between the density of 
damaging rock bursts as a function of the 
rate of energy release, based on mining 








- I 


<D 


I I 


I / 






sz 










.Q 


en 


03 


Severe 


Extreme / 




en 


™ 


^~ 






20 


2 


CO 


O 
2 






15 










/ — 


10 


- 








- 


5 


- 


^ 




I I 


I 



o 

CD *- 

<co 

^ O 
< *~ 
Q CC 

5* 

cc \- 

W CO 

CD CC 

^ => 

Z> CO 



2 4 6 8 10 

RATE OF ENERGY RELEASE, 
10 6 ftlbf/ft 2 

FIGURE 4.-lncrease in incidence and severity of rock 
bursts with increase in energy release rate. 

experience in the Republic of South 
Africa (38) , is shown in figure 4. The 
figure also shows the increase in rock 
burst severity experienced with increases 
in energy release. 

CONDITIONS FOR A SEISMIC EVENT 

In order for stored strain energy to be 
released as kinetic or seismic energy 
from a region of highly stressed rock, 
four conditions are necessary. The rock 
must be strained to the point of unstable 
equilibrium; there must be an additional 
increase in stress to trigger the rock 
failure; the nature of the failure must 
be a violent, brittle fracture; and there 
must be sufficient stored strain energy 
available to be released as kinetic 
energy. 

Unstable Equilibrium 

The condition of unstable equilibrium 
of a system is said to exist when the 
system is in equilibrium and its poten- 
tial energy is at a maximum value. If an 
attempt is then made to add any addition- 
al energy, the equilibrium will be upset 
and the potential energy of the system 
will be released as kinetic energy. The 
penedulum shown in figure 5 is a simple 
example of this concept. When the pendu- 
lum is rotated to the top dead-center 



14 




o 



KEY 
z Height 

Pendulum at unstable 

equilibrium 
,-, Pendulum at stable 
"-' equilibrium 

FIGURE 5. -Simple pendulum, illustrating unstable and 
stable equilibrium. 

position as shown, work is done on the 
system by raising the pendulum through 
height z, and energy is stored by the 
virtue of the pendulum's position. The 
system is then in equilibrium, and any 
attempt to do additional work on the 
system will upset the pendulum, causing 
it to fall and its potential energy to be 
released as kinetic energy. Once upset, 
the pendulum will swing back and forth 
through bottom dead center until the 
oscillations are dampened and the pendu- 
lum comes to rest in a position of stable 
equilibrium. As will be discussed later, 
these oscillations are analogous to seis- 
mic waves radiating from a seismic 
event. 

The condition of unstable equilibrium 
in a rock mass means that some point or 
local region has been strained by natural 
and induced stresses to the threshold of 
failure. If just the slightest amount of 
external work is then done on the rock in 
an attempt to strain it further, equilib- 
rium will be upset and failure will 
occur. The stored strain energy in the 
rock near the point of failure will then 
be released. If any of this energy 
should be released as kinetic energy, a 
seismic event, and possibly a rock burst, 
will occur. 



Triggering Stress Increase 

When the condition of unstable equilib- 
rium has been established, an additional 
stress increase must occur to trigger the 
rock failure. This stress increase, how- 
ever slight, may come from the zone of 
concentrated stress that advances in the 
rock ahead of excavation, or as is often 
the case, it may come from the sudden 
stress change caused by an excavation 
blast. 

The zone of concentrated stress induced 
in the rock ahead of an opening is shown 
in figure 6. The concentration of stress 
produces an inelastic, fractured zone 
around and ahead of the opening. If the 
rock is at or near unstable equilibrium 
ahead of the face, the approaching stress 
increase may initiate failure. Geologic 
discontinuities such as dikes or faults 
can cause an abrupt increase in the 
stress and trigger failure as the opening 
approaches the discontinuity. 

Violent Failure 

Once rock failure is initiated, it must 
be a violent, brittle type of failure and 
as such must be accompanied by an abrupt 
postfailure loss of strength. This con- 
dition requires a rock that is strong 
and hard, and exhibits brittle-material 
behavior upon failure. A stress-strain 
curve for a perfectly-brittle rock mate- 
rial is illustrated in figure 7. When 
such a material reaches its maximum 
strength and fails, the sudden loss of 
strength permits a violent displacement 
of fracture surfaces and a simultaneous 
drop in stress. Violent displacement of 
the fracture surfaces will cause the sur- 
faces to recoil and oscillate about their 
equilibrium position until the oscilla- 
tions are dampened. This motion is the 
mechanism responsible for conversion of 
the stored strain energy into kinetic, or 
seismic, energy. 

Release of Seismic Energy 

The violent displacement of fracture 
surfaces that releases and radiates seis- 
mic energy from the point of rock failure 
requires the availability of excess 



15 



Elastic zone 




FIGURE 6.-Mining-induced stress concentration ahead of face. 



1 


i 












"c 










i 
Str 


i 
ess 


CO 
CO 

LU 
CC 

r- 
CO 










dr 

< 


op 










KEY 




/ °c 


C 


o m 


pressive strength 



STRAIN 
FIGURE 7.-Stress-strain curve for perfectly-brittle material. 

strain energy. Excess energy means more 
energy than can be dissipated by stable, 
nonviolent rock failure and deformation. 
Energy availability is related to the 



stiffness of the strained rock around the 
fracture surfaces and the slope of the 
postfailure stress-strain curve. 

The stress-strain curve shown in figure 
8j4 illustrates stability following fail- 
ure because the slope of the postfailure 
curve is less than that, of the rock 
stiffness. In this case, there is no 
excess strain energy available to be 
released as kinetic energy. The stress- 
strain curve of the rock in figure SB 
illustrates a postfailure curve with the 
slope greater than that of the rock 
stiffness. Rock failure under these con- 
ditions is intrinsically unstable because 
increased displacement is accompanied by 
decreased strength or strain softening, 
and the surrounding rock will continue to 
load the fracture surfaces following 
initiation of failure. Continued loading 
will result in the violent displacement 
that releases excess strain energy as 
kinetic energy, until equilibrium is 
reached when the falling strength stabi- 
lizes at the fractured strength level. 



16 




to 
w 

LU 
DC 
I- 
to 



t s 






KEY 






a 1 


Solid 
strength 




4 


a 2 


Fractured 




strength 






4-4 


Rock 
stiffness 


18888 


Excess 








strain 








energy 




* 




^4 



STRAIN 

FIGURE 8.-Stress-strain curves. A, Nonviolent failure; 
B, violent failure due to excess strain energy. 



FAILURE MECHANISMS 

Rock, failure mechanisms involve loading 
conditions, physical properties of the 
rock, and material instability character- 
istics that combine to produce a fracture 
or failure zone and promote fracture 
growth once started. The understanding 
of failure mechanisms is an active and 
important area of current rock burst 
research. By establishing rock instabil- 
ity and failure criteria, and then 
incorporating these into numerical 
models, it may become possible to analyze 



mine geometries and calculate the proba- 
bility of rock bursting. 

Intact Rock 

The first efforts to establish failure 
criteria for brittle rock materials dealt 
with the initiation of failure and the 
development of failure planes under shear 
and normal stress conditions. The 
Coulomb criterion (35) postulates shear 
failure as a linear function of material 
cohesion and angle of internal friction 
across a failure plane, while the Mohr 
criterion postulates a functional rela- 
tionship, which is dependent on the mate- 
rial, between the shear and normal 
stresses. The Griffith criterion for 
brittle fracture postulates that failure 
is initiated with the growth and coal- 
escence of preexisting microf issures 
within the stressed rock. The microfis- 
sures ultimately join together to form 
the macroscopic failure surface. 

While these early efforts concentrated 
on failure initiation, later research 
included the behavior of rock in the 
postfailure portion of the stress-strain 
curve and the effects of confining stress 
and elevated temperature on rock response 
during loading and failure. 

Analytical models of postfailure behav- 
ior attempt to model failure localization 
and development within an initially uni- 
form material as an instability in the 
constitutive description of deformation. 
When stressed, the material undergoes a 
normal homogeneous deformation until a 
bifurcation point is reached. Continued 
loading then produces localization and 
development of a planar shear band at 
some point in the material where non- 
homogeneous deformation and eventual 
failure occur. Outside of the shear 
band, equilibrium is maintained and 
normal homogenous deformation continues 
with continued loading (39). 

Another model, which has application to 
bedded or layered rock, treats the rock 
failure mechanism as a problem in sur- 
face instability and may be considered a 
special case of the previously described 
model. This model develops an instabil- 
ity criterion in terras of the 



17 



uniaxial and compressive strengths, and 
applies a bifurcation analysis technique 
for failure localization (40). 

Faulted Rock 

The failure mechanism of geologic dis- 
continuities such as faults, joints, and 
bedding planes is dependent upon the 
shear and normal stresses acting on the 
surface and the frictional properties of 
the surfaces. Shear displacement of the 
surfaces may be prevented by normal 
stress, which provides a clamping action 
through the static coefficient of fric- 
tion of the surfaces. 

If mining-induced stresses should upset 
the balance of stress on the discontinu- 
ity by either increasing the shear stress 
or reducing the normal stress, then dis- 
placement along part of the discontinuity 
will occur. Whether displacement is 
stable or unstable will depend upon the 
stiffness of the rock and the slip 
response of the discontinuity. 

The model of a block on a flat surface, 
shown in figure 9, illustrates the sta- 
bility conditions of a discontinuity. 
The normal force N acts on the block 
perpendicular to the surface, while the 
tangential force T acts parallel to 
the surface through the spring having 
stiffness k. The system will be stable 
and the block stationary as long as 



y s N, 



(8) 



where u s is the static coefficient of 
friction and u s N is the static friction 
force. When the tangential force T is 
increased, or the normal force N is 
decreased, to the point that 



PsN, 



(9) 



the system will be in unstable equilibri- 
um. If there is then the slightest 
increase in force T, or reduction in 
force N, the block will begin to slide. 
Once sliding begins, the kinetic coeffi- 
cient of friction u k , being less than the 
static coefficient of friction, will 
allow the block to continue to slide with 
less force than that required to initiate 
sliding. 



-^mmw^ 



— t 



7 — 7 7 7 — 7 — 7 — 7~~7 — 7 — 7 — 7 — 7 — 7 — 7 — 7~ 

KEY 

k Spring stiffness 

N Normal force 

T Tangential force 

FIGURE 9.-Simple spring-loaded sliding block model. 

If the tangential force T required to 
extend the spring is greater than the 
force required to maintain sliding on the 
block -surf ace interface, the motion of 
the block will be stable. If, on the 
other hand, the force required to extend 
the spring is less than that required to 
overcome static friction, strain energy 
will be stored in the spring as the block 
is loaded by the tangential force T prior 
to motion. When the static friction is 
finally overcome and the block begins to 
move, the stored strain energy will be 
released on the block, causing an un- 
stable, violent motion. 

SOURCE LOCATIONS AND ROCK BURSTS 

Whether a rock burst will result from 
the release of seismic energy will depend 
on the magnitude of the energy released 
and the distance between an opening and 
the point of rock failure. When the 
radiated seismic energy wave arrives at 
an opening, the wave is reflected by the 
surface of the opening. If the magnitude 
of the seismic energy is sufficient, 
reflection of the seismic wave will frac- 
ture the surface rock and cause it to be 
violently expelled into the opening. 

At the Face 

If high stresses in the rock immediate- 
ly adjacent to an opening's fractured and 
inelastic zone should initiate a seismic 
event, the rock that fails will be 
expelled into the opening with explosive 



18 



force. Ro 
of seismi 
face are 
terms of 
damages, 
caused by 
are relati 
with othe 
proximity 
energy tha 
burst. 



ck bursts caused by the release 
c energy at or very near the 

known for their severity in 

fatalities, injuries, and 

Ironically, these bursts may be 

seismic energy releases that 
vely low in magnitude compared 
r seismic events. It is the 
of the opening to the source of 
t causes the severity of the 



Ahead of the Face 



Another cause of seismic activity 
responsible for rock bursting is the for- 
mation and propagation of shear fractures 
in the stressed rock ahead of an opening. 
That these are shear fractures is 
inferred from the detailed analyses of 



microseismic wave forms originating from 
ahead of openings (41 ). The locations of 
these shear failure planes and their 
orientations with respect to stopes and 
structural features such as dikes have 
also been inferred from microseismic data 
(42). As previously mentioned, features 
such as dikes may cause abrupt increases 
in the induced stress zone ahead of an 
advancing opening and thus promote seis- 
mic activity. 

In the bedded quartzites of the Repub- 
lic of South Africa's gold mines, shear- 
type fractures have been classified 
according to their inclination to the 
horizontal and their seismic activity 
(43). They have also been studied by 
excavating exploratory drifts and raises 
into the rock ahead of the face (44). 
Figure 10 illustrates the rock parting 






KEY 




P art ing planes 


Sh 


ear f racturin g : 


III 


Type 1 


\ 


Type 2 


^ 


T ype 3 



FIGURE 10. -Types of shear fracturing in bedded quartzite ahead of stope. (See text for 
discussion of types.) 



19 



planes and the three types of shear frac- 
ture commonly found. Type 1 fractures 
are, the most common, have a near-vertical 
dip, and are generally planar. Type 2 
fractures dip 60° to 75° and may form 
complex fracture zones as well as con- 
jugate fractures. Type 3 fractures dip 
20° to 40° and are usually found at or 
very near the surfaces of openings. 

As in the case of rock bursts from 
seismic events at or very near the face 
of an opening, bursts originating from 
shear fracturing often cause severe dam- 
ages because the seismic energy is 
released close to the opening. These 
fractures are often located within four 
or five opening widths of the face. 
Seismic energy magnitudes from these 
shear fractures have been measured in the 
range of 0.5 to as high as 3.0 on the 
Richter scale. 

On Faults 



Geologic discontinuities such as faults 
and joints are a third source of seismic 



activity and rock bursting in mines. As 
mentioned in the earlier section on fail- 
ure mechanisms in faulted rock, mining- 
induced stress changes may increase the 
shear stress or reduce the normal stress 
on a discontinuity and allow surface dis- 
placement. If the displacement is an 
unstable, violent motion, seismic energy 
will be released. 

The Richter magnitude of the seismic 
energy released by violent discontinuity 
displacements is often very large, in the 
range of 3.0 to 5.0, but the damages 
caused by rock bursts produced by these 
events is often less than that of bursts 
caused by other sources. The reason is 
the greater distance usually found be- 
tween mine openings and seismic event 
locations on the discontinuities. These 
events may be felt simultaneously at 
several mines in a district and yet cause 
relatively little or no damage at any of 
them. When such events do occur close to 
mine workings, the resulting rock burst 
damage is catastrophic. 



COEUR D'ALENE ROCK BURST RESEARCH 



Research into the causes of rock bursts 
in the Coeur d'Alene District and the 
development of control and preventative 
measures began in the early 1940' s when 
the potential seriousness of the problem 
was first recognized. Since then, the 
microseismic method has become a refined 
research tool for rock-mass monitoring 
and for studying rock failure mechanisms. 
Rock burst control measures such as 
destressing, i.e., reducing the level of 
stress in rock, and rock preconditioning 
have been developed, and the improvement 
of stoping methods to control rock bursts 
continues today. 

MICROSEISMIC METHOD 

Hard-rock miners have long known that 
the crackling and popping sounds made by 
rock in underground openings are related 
to the stability of the rock. This "rock 
talk," as miners refer to it, has been 
taken as a warning of potential rock 
failure. Any abrupt change in rock talk, 
especially any increase in noise rate or 



amplitude, would prompt miners to vacate 
the noisy area until the rock either 
failed or quieted. 

Early Development 

In 1938, the Bureau initiated research 
to determine if a relationship existed 
between the seismic velocity through a 
loaded support pillar and the state of 
stress in the pillar. During the experi- 
mental phase of this research, while 
using sensitive seismic equipment, it was 
discovered that rock under stress emits 
subaudible or microseismic signals in 
addition to audible sounds or rock talk. 
The generation of these microseismic sig- 
nals in stressed rock implied a relation- 
ship between the time rate of the signals 
and the state of stress in the rock (45). 
This relationship was soon verified dur- 
ing a seismic investigation of pillars in 
a deep copper mine in northern Michigan. 
During the course of the investigation, 
the ground being monitored became seis- 
mically noisy after an initial quiet 



20 



period, with the rate of noise increasing 
for about 15 min before terminating in a 
rock burst less than 50 ft from the point 
of observation (46). This unexpected 
event led to the hypothesis that micro- 
seismic activity, if it always preceded a 
rock burst, might be used to predict or 
warn of an imminent burst. With this 
event, development of the microseismic 
method as a rock burst research tool had 
its beginning. 

The first mention of an investigation 
into the rock burst problems of the Coeur 
d'Alene District came in October 1941, 
only 4 months after the district's first 
rock burst fatalities at the Sunshine 
Mine (47). There was optimism that the 
microseismic method, then in its early 
stages of development, could provide a 
solution for the problem. Unfortunately, 
no systematic method of rock burst pre- 
diction or warning was then available, 
and it was not until the mid-1960's that 
the microseismic method would be applied 
in the district when methods for seismic 
source location became available. 

Development of the microseismic method 
began with both laboratory and field 
testing. The laboratory testing of rock 
specimens stressed by testing machines 
revealed that the microseismic noise rate 
increased with increasing stress and 
that, just prior to specimen failure, 
there was a dramatic increase in the 
noise rate (48). The source of both the 
audible and subaudible noise is the 
microf racturing and microdisplaceraents 
that occur as the rock responds to the 
increasing load. The dramatic noise 
increase as failure approaches comes from 
the increase in microf racturing and the 
growth of fracture surfaces on which 
failure occurs. 

Field testing with microseismic equip- 
ment was done by monitoring stressed 
pillars and rock structures in burst- 
prone mines (49). The object of this 
work was to monitor microseismic activity 
for a period and determine whether 
repeated patterns of noise increase 
preceded rock failure, and whether or not 
these patterns could be recognized. The 
equipment available at the time did not 
permit location of the noise sources, and 



consequently, the monitored noise in- 
cluded extraneous noise and was too 
irregular to provide recognizable pat- 
terns prior to failure. Also, the moni- 
toring method was highly subjective and 
heavily dependent upon the skill and 
experience of the user, which always left 
an uncertainty in the results. 

Seismic Source Location 

Research with the microseismic method 
as a means of determining rock stability 
continued with fieldwork in underground 
zinc, iron, copper, and oil shale mines 
in the United States (50), although few, 
if any, significant changes were made in 
the method until the mid-1960's. It was 
then, following the work of Cook ( 26 ) in 
the Republic of South Africa, that the 
Bureau began research with three-dimen- 
sional underground seismometer arrays for 
seismic source location. By this time, 
the quality and capabilities of available 
equipment had improved to the point where 
multiple-channel microseismic monitoring 
was possible and data could be recorded 
on magnetic tape. With these improve- 
ments came development of a broadband 
microseismic system capable of recording 
the true waveform of an event for 
detailed analysis ( 51 ) and multiple- 
channel systems to permit mine-wide 
determination of noise source locations 
(52). 

Rock burst research in the Coeur 
d'Alene District resumed with the appli- 
cation of noise source location tech- 
niques that were applied to the problem 
of pinpointing zones of high stress con- 
centration and, consequently, zones where 
a seismic event would most probably 
originate (53). The technique was tested 
in 1968 at the Galena Mine by locating 
the sources of microseismic activity 
during a 52-day period and preparing the 
cumulative noise source location plots 
shown in figure 11. These plots estab- 
lish the noise sources, delineate the 
high stress concentration zones by the 
contoured areas of dense microseismic 
activity, and also indicate the sources 
of rock bursts that eventually occurred 
in the study area. 



21 



• • 





14 days 



34 days 



Rock burst 
Day 34 




• • 




Rock burst 
Day 41 



3 7 days 



4 1 days 




Rock burst 
Day 42 



KEY 

• Noise location 

CJ3SJ) Noise density area 
k ■'■■■':■ ■■■■ : .i Mined and filled vein 



52 days 



FIGURE 11. -Cumulative plots of noise source locations. 



22 



Rock Burst Monitoring Systems 

Continued research by the Bureau in the 
early 1970's led to the development of a 
microseismic rock burst monitoring (RBM) 
system, followed by development of 
another monitoring system that uses a 
minicomputer (54). The original RBM sys- 
tem consists of an array of geophones 
with preamplifiers, high-gain amplifiers, 
the RBM control unit, and a printer. 
During operation, if a seismic event of 
greater than a preset threshold magnitude 
is detected by four or more geophones 
within a 0. 2-s interval, the first 
arrival times of the signals are 
recorded, the relative energy of the sig- 
nals is calculated, and the data are 
printed. The event location is then cal- 
culated from the first arrival times. 
The minicomputer RBM system also uses an 
automated data-gathering system but 
includes a minicomputer for automatic 
processing of the data, including calcu- 
lation of the event coordinates. 

One of the first district mines to 
install its own minicomputer RBM system 
was the Lucky Friday Mine, in 1972. The 
system monitored up to 23 geophones stra- 
tegically located throughout the mine, 
processed microseismic noise data, and 
plotted the noise location coordinates 
(55). The original RBM system installed 
at the Lucky Friday Mine has been 
replaced to take advantage of the more 
sophisticated equipment developed since 
the first system was installed. During 
the same period, other district mines 
such as the Sunshine, Galena, Crescent, 
and Star also installed RBM systems. 

The information obtained from micro- 
seismic monitoring systems is routinely 
used by mine personnel to detect and 
locate areas of high stress concentra- 
tions, after which control measures such 
as blasting to relieve stress or differ- 
ent stoping sequences to reduce mining- 
induced stresses are implemented. That 
these techniques have been successful 
in controlling rock bursts is indi- 
cated by a 1973-75 study of district 
mines (56). The results showed an over- 
all decrease in the ratio of damaging 
rock bursts to total bursts, where total 



bursts included those bursts triggered by 
destress blasting. The number of damag- 
ing rock bursts remained fairly constant 
over the study period, although the mines 
were being developed to deeper and more 
burst-prone levels. 

Rock Failure Mechanisms 



In addition to the use of microseismic 
systems at mine sites for rock stability 
and rock burst monitoring, such systems 
have been used in the laboratory and 
field to study rock failure mechanisms. 
Laboratory research during the late 
1960's used microseismic emissions to 
locate microf racturing in rock specimens 
under increasing load, which provided 
insight into microf racturing and failure 
plane development (57-58). 

The detailed analysis of microseismic 
field data provided valuable information 
concerning the influence of geologic 
structures on rock bursting, as well as 
the actual failure mechanisms. In many 
Coeur d'Alene District mines, noise 
source locations occur in bands that 
correspond to known structural features 
such as faults or hard quartzite layers 
interbedded with softer argillites. As 
previously mentioned, such structural 
controls are important because they may 
exert a significant influence on the con- 
centration of mining-induced stresses. 

DESTRESSING AND PRECONDITIONING 

Destress blasting and rock precondi- 
tioning are rock burst control techniques 
developed to reduce the potential for 
rock bursts by uniformly fracturing a 
large section of solid rock. While 
destressing is used to reduce stress con- 
centrations in highly stressed pillars 
remaining after a block of vein is mined, 
rock preconditioning is applied to a 
block of vein prior to mining and subse- 
quent buildup of induced stress. In both 
cases, fracturing is achieved by drilling 
blastholes into the vein, loading them 
with explosives, and firing. In prac- 
tice, a series of long, closely spaced 
holes are drilled into the vein from a 
crosscut, drift, or stope. The actual 



23 



hole pattern will depend to some degree 
on the available access to the vein and 
the drilling equipment used, but the 
ideal is a uniform distribution of holes. 

The object of both control techniques 
is to produce uniform fracturing of the 
vein being destressed or preconditioned. 
In the case of destressing, fracturing 
has the effect of relieving concentrated 
stress by allowing rock displacement on 
the newly formed fracture surfaces and, 
consequently, allowing the dissipation of 
stored strain energy under relatively 
controlled conditions. Stresses that 
were concentrated on the small area of 
the pillar are then redistributed over 
the much larger area of the surrounding 
filled vein. 

Since rock preconditioning is used on a 
block of vein prior to mining, fracturing 
of the rock has the effect of reducing 
its elastic modulus, and therefore its 
stiffness, before any abrupt increase in 
mining-induced stress. The rock burst 
potential is reduced because when the 
block is mined the rock is able to yield 
as stresses increase, rather than remain- 
ing rigid until increasing stresses cause 
a violent failure. 

Pillar Destressing 

The pillar destressing technique was 
first used in the Coeur d'Alene District 
during the late 1960's at the Galena 



Mine (31 )♦ The purpose was to test the 
hypothesis that in order to produce a 
rock burst in a pillar, the stiffness of 
the pillar rock must be greater than the 
stiffness of the wall rock that comprises 
the loading system. When this condition 
exists and failure occurs, the stored 
strain energy in the wall rock will con- 
tinue to load the pillar following fail- 
ure, and kinetic energy will be released, 
producing a violent rock burst. 

Microseismic monitoring of a pillar at 
the Galena's 3700 level indicated devel- 
opment of a potential burst zone in an 
area that had recently experienced a 
burst. The stiffness of the pillar and 
wall rock were calculated from a numeri- 
cal model, and the pillar stiffness was 
found to be over twice that of the adja- 
cent wall rock. Under these conditions 
of increasing stress in a stiff pillar, 
the potential for another rock burst was 
considered great. 

In order to test the stiffness hypothe- 
sis, the pillar was destressed to reduce 
both its stiffness and the level of 
stress. Pillar and destress holes are 
shown in figure 12. The effectiveness of 
the technique was evaluated by the dif- 
ference in seismic velocities through 
the pillar taken before and after blast- 
ing. Since seismic velocity is propor- 
tional to the rock's effective elastic 
modulus and thus to its stiffness, 
a decrease in velocity would imply a 



,\VF7fc_ 






^ 



■1 .1 



J 



jm^ 



\\\HL 



M& 



3,700-ft level 




Destress holes 



Sand- filled 



10 20 30 




Section A- A 



Scale, ft 
FIGURE 12. -Holes drilled into pillar from stope below for destress blasting. 



24 




A1&& 



BEFORE DESTRESSING 
JMA ■ — V&U. 



KEY 

• Geophones 
o Seismic velocity 
shot hole 

I ) Seismic velocity, ft/s 




Sand-filled 

13 5 raise 



AFTER DESTRESSING 
FIGURE 13. -Seismic velocity contours as measured before and after destress blasting. 



reduction in stiffness. As shown by the 
velocity contours in figure 13, blasting 
did result in an overall decrease in 
seismic velocity. In addition, calcu- 
lated pillar stiffness following blasting 
was less than that of the wall rock. The 
objective of destressing had been 
achieved, and the pillar was subsequently 
mined without further incident. 

Rock Preconditioning 

Based on the earlier success with the 
destressing technique, the idea of pre- 
fracturing or preconditioning a block of 
vein prior to mining was developed and 
demonstrated in a district mine. 



First Demonstration 

The first use of the rock precondition- 
ing technique in the Coeur d'Alene Dis- 
trict was in 1976 at the Hecla Mining 
Co. 's Star Mine (59) . During development 
and mining of the 7500 level, unusually 
heavy rock bursting was encountered. 
After studying the stoping method being 
used, the stope geometry, geology, and 
stress conditions, Bureau researchers and 
mine personnel concluded that heavy rock 
bursting would also be encountered while 
mining the levels below the burst-prone 
section of the 7500 level. 



25 



The decision was made to test rock pre- 
conditioning by demonstrating its use on 
a small section of the 7700 level prior 
to mining. The block of vein selected, 
as shown in figure 14, was located 
directly below the area of heavy bursting 
encountered on the 7500 level and was 
accessible by the 7 and 10 crosscuts. 
Experience has shown that there are two 
times during the mining cycle when burst- 
ing is most probable: during opening of 
the initial or "I" drift at the bottom of 
a vein block and while mining the pillar 
produced at the top of the vein block. 
Preconditioning the vein 40 ft above and 
below the 7700 level would allow mining 
up from the 7700 level to begin in pre- 
conditioned rock and would also allow 



later mining of the 7900 level to end in 
a preconditioned pillar. 

Since mining of the initial drift from 
the 7 crosscut had already begun, the 
holes from this crosscut could not be 
drilled into the vein as were those from 
the 10 crosscut, but had to be located 
about 20 ft from the vein (fig. 15). The 
quartzite wall rock, not the vein, was 
preconditioned from the 7 crosscut. The 
blastholes were drilled and loaded from 
the crosscuts, as shown in figure 16, and 
detonated. 

The preconditioning was monitored and 
evaluated by means of seismic velocity, 
stress, and displacement measurements. 
The reduction in seismic velocity fol- 
lowing blasting indicated significant 




= = = = 6 ,9 00-f t level 



7 , 1 00-ft level 



7 ,300-f t level 



7, 5 0-ft level 



7 ,700-f t level 



= = =7,900-ft leve 



Scale, ft 



FIGURE 14.-Vein preconditioning, first demonstration site, 7,700-ft level of Star Mine. (Numbers 
identify crosscuts.) 



26 



LEGEND 

1/ Crosscut 

PZ^ Blasted zone 

Raise development 
lateral 




50 100 

I I I 

Scale, ft 



1 03 



FIGURE 15.-Plan view of preconditioned zones, 7,700-ft level haulage lateral, and crosscuts to vein. 



fracturing of the vein around the 10 
crosscut and the wall rock around the 7 
crosscut. Stress gauges, installed in 
the wall rock prior to blasting, showed a 
reduction in the horizontal component of 
stress, and displacement measurements 
showed an induced closure across the 
vein. 

The real test came during subsequent 
mining. Since the wall rock, and not the 
vein, had been prefractured at the 7 
crosscut, popping rock and typically dif- 
ficult conditions were encountered when 
work resumed on the initial drift. Stop- 
ing of the first four floors, or 40 ft, 



was accompanied by increased microseismic 
activity and one major rock burst. 
Stoping from the 10 crosscut, in prefrac- 
tured vein, was without the usual rock 
popping or seismic activity. Even the 
initial drift was completed without prob- 
lems, and during raining of the first 40 
ft, microseismic activity was much lower 
than that from the 7 crosscut. During 
continued stoping of the 7700 level, 
above the influence of the preconditioned 
zone, seismic activity and rock bursting 
resumed, as was expected to happen with- 
out preconditioning. 



27 



Pounds of explosives 

i 



crosscut 



a 




crosscut 



o^ 



10 20 



Scale, ft 
FIGURE 16. -Preconditioning drill-hole patterns and loading, 7 and 10 crosscuts, 7,700-ft level. 



28 



Second Demonstration 

The experience gained by precondition- 
ing a small section of the Star Mine's 
7700 level and the reappearance of rock 
bursting above this section indicated the 
need to test the technique on an entire 
stope block. The location of this large- 
scale second demonstration was the Star 
Mine's 7900 level, as shown in figure 17. 
This area was chosen because it included 
the previously preconditioned zone above 
and was directly below the area of heavy 
bursting encountered on the 7500 and 7700 
levels (60). 

A primary safety concern about precon- 
ditioning the entire stope block was that 
prior horizontal stress on the 7900 level 
would be transferred by the precondition- 
ing to joillars between the 7500 and 7700 
levels. Any such increase in pillar 
stress would increase the pillars' rock 



burst potential dramatically. To deal 
with this problem, preconditioning was 
divided into two phases separated by 
pillar destressing from the 7500 level 
and initial mining on the 7900 level. 

The first phase, which was completed in 
1979, consisted of preconditioning a 
narrow vein section from the 8 and 12 
crosscuts on the 7900 level, while mining 
continued on the 7700 level above. The 
vein was drilled and loaded, as shown in 
figure 18, to produce a preconditioned 
area extending 50 ft above and below the 
level. Following preconditioning, mining 
from the 8 and 12 crosscuts began in the 
prefractured rock with little microseis- 
mic activity and no rock bursting. 

Before the remaining stope block could 
be preconditioned in the second phase, 
the pillars between the 7500 and 7700 
levels had to be destress blasted. 
Drilling was begun from the 9 crosscut on 




ite ' / 

Axis of burst zone ° 10 ° 20 ° 

I _l I 

Scale, ft 



= 6,900-ft level 



= 7, 1 00-ft level 



= 7,300-ft level 



= 7,500-ft level 



===== = = 7,700-ft level 



7,900-ft level 



=z = = 8.100-ft level 



FIGURE 17. -Vein preconditioning, second demonstration site, 7,900-ft level of Star Mine. (Numbers identify 
crosscuts.) 



29 



Pounds of explosives 





Scale, ft 
FIGURE 18.-Preconditioning drill-hole patterns and loading, 8 and 12 crosscuts, 7,900-ft 



level. 



30 




Scale, ft 



FIGURE 19.— Pillar destress blastholes drilled from 9 crosscut, 7,500-ft level. 



7500 level, as shown in figure 19, but 
before drilling could be completed, a 
major rock burst occurred. The burst, of 
magnitude 2.6 on the Richter scale, 
effectively relieved stresses in the pil- 
lar below the 9 crosscut. The pillar 
below the adjacent 5 crosscut was drilled 
and blasted without any problems. 

The second phase was completed in 1980 
and consisted of preconditioning the 
remaining stope block by drilling ver- 
tically from the by-then-corapleted second 
floor of the 8 and 12 stopes. The drill 
holes are shown in figure 20 with the 
first phase area on the 7900 level. Fol- 
lowing blasting of the stope, mining 
resumed with only moderate microseisraic 
activity and no rock bursting until near 



the top of the 8 stope. Microseismic 
activity in this area increased until a 
rock burst of 1.9 on the Richter scale 
occurred in a vein that had not been 
adequately pref ractured. The vein was 
further destressed by drilling down from 
the 7 crosscut above and blasting (_8). 

While destressing and preconditioning 
have been shown to be effective rock 
burst control measures, their use 
involves some risk and high costs. When 
a block of vein is fractured by either 
method, some of the stress on the block 
becomes redistributed in the adjacent 
intact rock. This increases the level of 
stress on the intact rock and conse- 
quently increases the probability of a 
seismic event and resultant rock burst. 



31 




7,700-ft level 



7,900-ft level 



FIGURE 20. -Preconditioning hole pattern, drilled from 7,900-ft-level stope. (Numbers identify crosscuts.) 



The techniques must therefore be used 
with careful planning and very close 
monitoring. 

The major disadvantage of these stress 
control techniques is their high cost. 
Both techniques are labor- and equipment- 
intensive since a great deal of blasthole 
drilling must be done. To this, the cost 
of explosives must be added. These 
expenses only add to the already high 
costs of raining each ton of ore. 

STOPING METHODS AND MINE GEOMETRY 

The influence that a particular mining 
method will have on rock bursting at 
depth is seldom a consideration during 
the initial planning and development of a 
mine. Historically, underground mining 
choices have been based upon ore deposit 



and host rock properties, geologic struc- 
ture, mine production requirements, and 
economic factors. The relationship 
between a mining method and rock bursting 
potential does not become a problem until 
bursting becomes a problem. By that 
time, unfortunately, the mine has usually 
been developed to a considerable depth 
and the raining method well established 
and difficult to change or modify. 

Overhand Stoping 

Coeur d'Alene District mines have com- 
monly used the overhand cut-and-fill 
stoping method. This method is typically 
used to mine steep, narrow, high-grade 
vein deposits in strong, competent wall 
rock. The vein is divided into levels, 
usually 200 ft apart vertically, and each 



32 



level is mined from bottom to top, hence 
the name overhand. A vertical section of 
several mined and filled levels is shown 
in figure 17. 

Access to the vein on each level is by 
crosscuts driven from a lateral haulage 
drift paralleling the vein, as illus- 
trated in figure 15. The crosscuts, 
located about 200 ft apart horizontally, 
divide the vein into blocks. Mining pro- 
ceeds from the crosscuts by "raising up" 
(excavating upward) into the vein and 
then excavating a series of horizontal 
stopes. Mining is done on a drill, 
blast, and muck cycle, and as each stope 
is mined out, it is backfilled with sand 
or waste material. Timber supports are 
added in some areas. The fill material 
provides support for the wall rock as 
well as a surface from which to continue 
mining above. Figure 21A shows a typical 
overhand stope with its raise and ore 
chutes, backfill, and active stope. 



As each stope is mined and filled, the 
horizontal stresses become redistributed 
in the remaining intact vein above and 
below. Initial wall-rock convergence 
compacts the fill, and resistance of the 
fill to further compaction provides wall- 
rock support. While the fill modulus is 
normally much less than the wall-rock 
modulus, possibly as little as one 
one-hundredth, the support provided has 
been shown to be significant in terms of 
the stress induced in the remaining vein 
(61). The fill then provides more than 
just passive wall-rock support and a sur- 
face from which to continue mining — it 
plays an active role in the control of 
mining-induced stresses. 

Overhand stoping produces a pillar of 
vein rock as mining proceeds upward 
toward the filled level. Typical pillars 
are shown in figures 17 and 19. The 
redistribution of stress with continued 
mining results in an increased stress 





FIGURE 21. -Generalized drawings of {A) overhand and (B) underhand stoping methods. 



33 



concentration in the pillar until 
relieved by pillar failure. The nature 
of the pillar failure will be violent or 
nonviolent depending upon the relative 
stiffnesses of the pillar and wall rock. 
A nonviolent failure is characterized by 
gradual fracturing and yielding, while 
violent failure is often characterized by 
the release of seismic energy and a rock 
burst. Pillar bursts in overhand stopes 
account for many of the rock bursts 
experienced in the Coeur d'Alene District 
and are often responsible for severe 
damages since they are located directly 
above the active stope. 

As district mines have been developed 
to greater depths and, consequently, into 
more highly stressed ground, rock burst- 
ing has increased in both incidence and 
severity. One approach to the allevia- 
tion of this problem has been the inves- 
tigation of alternative stoping methods 
that minimize or reduce induced stress 
concentrations. Research by the Bureau 
has led to the development and testing of 
underhand cut-and-fill stoping methods as 
a means of reducing the rock burst hazard 
by eliminating the pillar in highly 
stressed narrow veins. 

Underhand Stoping 

Underhand stoping methods typically 
consist of mining a block of ore by cut- 
ting and filling in sequence from the top 
of the block to the bottom. A general- 
ized drawing of a section of vein being 
mined by the underhand method is shown in 
figure 21b. Mining is done on a drill, 
blast, and muck cycle as with the over- 
hand method, but downward since the 
intact vein is below the stope level. 
Following each cut, the open stope is 
backfilled with a cement-stabilized fill 
material. A cemented fill is necessary 
to prevent the fill from caving into the 
stope below when it is opened. Addition- 
al support and reinforcement techniques 
such as wire mesh, timbers, and rock 
bolts may also be required to stabilize 
the lower portion of each fill layer. 



The primary advantage of the underhand 
stoping method over the overhand method 
with respect to rock bursting is the 
absence of a pillar. Since the mining 
direction is downward into a continuous 
vein, a highly stressed pillar is not 
formed. Instead, abutment stresses 
remain in the intact vein and wall rock 
immediately below the open stope and move 
downward as mining proceeds downward. 
Above the open stope, the cemented fill 
is compacted by wall-rock convergence, 
and displacement of the wall rock pro- 
duces a region of reduced stress. 

If, for some reason, the underhand 
method is simultaneously used on more 
than one level, sill pillars will be 
formed as mining approaches the bottom of 
each level. These pillars will then 
undergo the same increasing stress con- 
centrations as pillars formed by the 
overhand method. A model for simulating 
various overhand and underhand stoping 
sequences is shown in figure 22 along 
with the order of elements mined. The 
calculated energy release rates for these 
simulations are shown in figure 23a 
and B. Figure 23a compares energy 
release rates of overhand and underhand 
stoping methods, and illustrates the 
abrupt increase in energy release rate as 
the overhand pillar is mined. The under- 
hand method results in a uniform rate of 
energy release. Figure 23b compares 
energy release rates for single- and 
multiple-level underhand stoping se- 
quences. The energy release rates pro- 
duced by multiple-level underhand stoping 
approach those of overhand stoping when 
sill pillars are formed and mined (62). 

While multiple-level underhand stoping 
presents an increased risk of rock burst- 
ing, the method still retains advantages 
over the overhand method. One advantage 
is that if pillar destressing or rock 
preconditioning techniques are applied to 
the vein, subsequent mining will proceed 
from above the prefractured rock and not 
from below it. This eliminates the haz- 
ard associated with working under frac- 
tured and possibly loose rock. Another 



34 



Mined 


and 






Level 1 


i 
-> 


1 

2 




3 




4 


c. 


5 


o 


6 




i 


7 




8 




9 




10 


I _ ,, „ I o 




Level d. , 


i 


1 1 




12 




13 




14 


c 


1 5 


o 


16 






1 7 




18 




1 9 




20 






Level o — 




I 



filled 



Simulation 
Overhand stoping 

Underhand stoping 

Multilevel underhand 
stoping (40-ft pillar) 

Multilevel underhand 
stoping (60-ft pillar) 

Multilevel underhand 
stoping (100-ft pillar) 



10' 



Vein 



Order of elements mined 

10, 9, 8, 7, 6, 5 and 20, 4 and 19, 
3 and 18, 2 and 17, 1 and 16 

1, 2, 3, 4, ... 20 

1, 2, 3, 4, 5, 6, 7, 8, 9 and 1 1, 
10 and 12, 13, 14, 15, 16, 17, 18, 
19, 20 

1, 2, 3, 4, 5, 6, 7, 8 and 11, 9 and 12, 
10 and 13, 14, 15, 16, 17, 18, 19, 20 

1, 2, 3, 4, 5, 6 and 1 1, 7 and 12, 

8 and 13, 9 and 14, 15, 16, 17, 18, 19, 

20 



FIGURE 22.-Model for simulated overhand and underhand stoping of section of vein. 



advantage is that if a rock burst should 
occur, it will be more apt to occur in 
the vein below the stope. The stope 
damage would probably be less severe, 
with safer cleanup and repair work, than 
if a rock burst above caused caving into 
the stope. 

Underhand Stoping Demonstration 

Initial research into underhand stoping 
methods as a means to reduce the rock 
burst potential in Coeur d'Alene District 
mines was conducted in two phases. Dur- 
ing the first phase, several mines in the 



United States and Canada that use the 
method were examined and studied to 
determine current practices. Feasibility 
and cost-effectiveness assessments were 
made, and based upon the study results, 
initial design recommendations were 
prepared. 

The second phase of research consisted 
of a demonstration test of underhand cut- 
and-fill mining in a district mine. The 
demonstration site had to have a history 
of rock bursting, include pillar destress 
blasting, and be in an active mining 
location so that a cost comparison could 
be made. 



35 



CM 
C 5 



A KEY 

Underhand stoping 

Overhand stoping 
Multilevel underhand 
stoping: 

40-ft pillar 

60-ft pillar 

1 OO-ft pillar 




20 40 60 80 100 

EXTRACTION, pet 

FIGURE 23. -Energy release rates for simulated stoping. 
A, Overhand and underhand stoping; B, multilevel underhand 
stoping. 

The site chosen was the 6700 to 6900 
level of the Grouse vein at the Hecla 
Mining Co. 's Star Mine. This portion of 
the vein, shown in figure 24, had been 
mined by overhand stoping to within 50 ft 
of the 6700 level. A rock burst had 
occurred in the pillar near the 125 



raise, and based on the vein's history, 
additional bursts were anticipated as the 
remaining pillar was mined. 

The plan for the underhand stoping 
demonstration was to destress-blast the 
remaining pillar by drilling vertically 
from the stope below, backfill the 6700- 
level haulage drift, and mine the pillar 
from above by underhand cut-and-fill 
stoping. 

To monitor the demonstration, instru- 
ments were installed in the 6700-level 
haulage drift. Included were extensome- 
ters for wall closure, stress gauges for 
wall-rock stress changes, and soil pres- 
sure cells for measuring loading on the 
cemented fill. Extensometers and pres- 
sure cells were added to each stope while 
that stope was open to continue monitor- 
ing closure and fill loading as the pil- 
lar was mined. The seismic velocity 
through the pillar was determined both 
before and after destress blasting to 
check the effectiveness of fracturing. 

While instruments were being installed, 
the 6700-level haulage drift was prepared 
for the cemented fill by construction of 
a floormat to restrain any loose fill 
during subsequent mining below. The 
floormat construction details are shown 
in figure 25. The wire mesh and porous 
fabric were necessary to retain solids 
while draining water from the cemented 
fill. As part of the demonstration, an 
experimental cement plant was set up for 
preparing and pumping fill material from 
the surface to the stope. 

Destress blastholes were drilled ver- 
tically into the pillar, except in the 
future raise areas shown in figure 26. 
These areas were not blasted because it 
was feared that fractured rock would 
cause problems when the raises were 
driven. Once the holes had been drilled 
and loaded and the 6700-level drift 
filled, the pillar was blasted. 



36 




6,700-ft level 



6,900-ft level 



FIGURE 24.-Underhand stoping demonstration site, 6,700-ft level of Grouse vein, Star Mine. (Numbers 
identify crosscuts.) 



Raise driving up through the pillar to 
the 6700 level began at the 122, 125, and 
129 stopes following blasting. The 122- 
stope raise was driven without diffi- 
culty, but some problems were experienced 
with the 125- and 129-stope raises. The 
125 raise passed through the region of 
rock fractured by a prior rock burst, and 
loose, broken rock became a problem while 
the raises were driven up through the 
pillar. Problems on the 129 raise were 
related to overloaded blasting rounds, 
which caused excessive fracturing of the 
already fractured rock. 



Following completion of raise driving, 
the first underhand stoping cut was begun 
beneath the cemented fill of the 6700 
level. As mining proceeded in each di- 
rection from each raise, two problems 
became evident. The first was that the 
quality of the cemented fill was incon- 
sistent, and the second was that fill 
material in the drift above crushed and 
buckled because of the large amount of 
wall-rock closure. 

The quality problems were the result of 
inconsistent cement and fill mixing and 
poor proportioning of the fill material, 



37 




12- by 12-in stringer 
FIGURE 25.-Floormat construction details for cemented fill material. 



which caused a separation and loss of 
cement from the fill during pouring. The 
cement and fill mixing problem was easily 
corrected at the surface mixing site, but 
the proportioning of fill material con- 
tinued to be a problem. During the fil- 
ling of later cuts, the problem was par- 
tially corrected by filling the first 3 
or 4 ft and then allowing the cement to 
settle before adding the remaining fill. 

The great amount of wall-rock closure 
following mining indicated that pillar 
destress blasting had been successful but 
caused problems in maintaining the raise 
openings. As a consequence, continued 
raise repair work was required while the 
pillar was mined. Extensometers in the 



6700 level indicated about 4 in of clo- 
sure with the first cut, but during the 
year required to mine the pillar, approx- 
imately 14 in of additional closure 
occurred in the raise areas. 

Four underhand cuts were required to 
mine the pillar. The only problems 
encountered during mining involved over- 
loaded blastholes and blastholes drilled 
too close to the filled stope above. 
These damaged the floormat and fill 
above, but were easily corrected with 
experience and more careful drilling and 
blasting procedures. 

Overall, the project successfully 
demonstrated that underhand cut-and-fill 
stoping with massive destress blasting 



38 




FIGURE 26.— Pillar destress blastholes drilled from stope below 6,700-ft level. (Numbers identify crosscuts.) 



can be used in the Coeur d'Alene District 
mines to control rock bursts. Even 
though the pillar was highly burst prone, 
only one moderate burst took place during 
the project. That burst was located in 
the wall rock and occurred during the 
first cut. Damage from the burst was 
minor and limited to some loose rock 
being shaken down in the nearby stopes. 

Lucky Friday Underhand Longwall 

Conceptual planning for the second 
demonstration of underhand cut-and-fill 
mining in the Coeur d'Alene District 
began in 1983 at the Hecla Mining Co. 's 
Lucky Friday Mine. The threat of in- 
creased rock bursting with continued min- 
ing at greater depths prompted a decision 
to develop an inherently safer mining 
method. In addition to providing a 



measure of control over rock bursting, 
the method would also have to provide the 
required production to be competitive. 

A site for the Lucky Friday underhand 
longwall was set aside between the 5100 
and 5300 levels and the 106 and 110 
crosscuts (fig. 27). This site provides 
a 500-ft length of narrow, nearly verti- 
cal vein at a point where the vein turns 
from a north-south to a northeast- 
southwest strike. The vein block above 
the 106 crosscut would be mined by the 
conventional overhand cut-and-fill method 
while at the same time the vein below 
would be mined by the underhand method. 

Initial plans considered the use of 
steel-lined mill hole raises in the vein 
as an ore pass from the stopes to the 
5300 level. Development and testing of 
the design and the experience with raises 
at the Star Mine led to the conclusion 



39 




ined and 
filled 



5,100-ft 
level 



5,300-ft level 



FIGURE 27. -Isometric view of Lucky Friday underhand longwall site. 



that boring and maintaining openings in 
the vein would present very difficult 
problems. In the existing high horizon- 
tal stress field, rapid deformation of 
the hole following boring and reaming 
could make it impossible to maintain a 
circular opening. Once completed, the 
risk of wall caving or spalling prior to 
or during liner installation would be 
high and would present a serious safety 
hazard. Furthermore, during subsequent 
underhand stoping, the expected wall-rock 
closure would aggravate any existing 
problems, making maintenance very 
difficult. 

Because of these potential problems 
with mill hole raises, a ramp system with 



an ore pass distant from the vein was 
developed, as shown in figures 27 and 28. 
The ramp allowed placement of the ore 
pass away from the area of high induced 
stress at the vein and also permitted the 
use of diesel-powered load-haul--dump 
units. During underhand stoping, ramp 
development was kept just ahead of mining 
to provide for vein access and ore haul- 
age. The ramp design consisted of alter- 
nating south, center, and north ramps, 
which spaced the openings farther apart 
for greater stability in the highly 
stressed wall rock. 

A part of the project requiring much 
research and development was that of the 
cemented-f ill system for backfilling the 



40 



Cutoff fault 

1 10 crosscut 




Ore pass 



,No. 2 shaft 




5, 100-f t level 
Floor 1 
Floor 2 



Access ramp 
North ramp 



South ramp 



Ore pass 



5,300-ft level 



FIGURE 28.-Plan and cross section of Lucky Friday underhand longwall site. 



stopes. Initial research centered on 
methods of preparing a fill material of 
adequate strength to resist crushing by 
wall closure. Closely coupled with the 
problem of preparation was the problem of 
transporting the fill mixture under- 
ground, since the transport system 
influences the equipment required for 
mixing. The fill system developed mixed 
mill tailings with cement in a concrete 
batch plant at the surface and then used 
a pump and a pipeline to transport the 
cemented fill to the stopes. 



A two-dimensional, finite-element model 
of a vertical section through the vein 
and ramp system was prepared to simulate 
both overhand stoping from the 5100 to 
the 4900 level and simultaneous underhand 
stoping from the 5100 to the 5300 level. 
The model simulated cut-and-fill stoping 
with sandfill for overhand stoping, ce- 
mented fill for underhand stoping, and 
development of the ramp haulage system. 
Analyses of the modeling results show 
the development of induced stresses 



41 



4,900 




5,100- 



Q. 
LU 
G 



5,300 



3 






KEY 



Lateral 
Ramp system 
Sand till 
' "■'■■' Cemented sand till 



<=> Vein 

^^ m Active mining area 

<3^ Stress concentration 



FIGURE 29. -Mining-induced stress concentrations from simulated overhand (above 5,100-ft level) and underhand 
(below 5,100-ft level) stoping. A, Start of stoping; B, half of each level completed; C, stoping nearly 
complete. 



with each method. The stress contours 
shown in figure 294, B, and C illustrate 
the concentration of stress in the pillar 
as overhand stoping approaches the 4900 
level. Stresses induced by underhand 
stoping generally remained in the vein 
and wall rock, below the active stope, 
while the ramp area remained clear of the 
high induced stresses near the vein. 

As in prior demonstrations, instruments 
were installed to monitor stresses and 
stope closure during mining. These 
instruments included extensometers in the 
ramp and stope openings and pressure 
cells in the filled stopes to monitor 
wall loading on the fill. In situ stress 
measurements were also taken, and all 
field data were used to verify the 



finite-element model as well as monitor 
the response to mining. 

Development of the ramp system to ap- 
proach the vein began in early 1985 and 
was followed by the start of underhand 
stoping at the 5100 level. When the 
first stope had been completed, it was 
backfilled with cemented fill, and the 
second underhand stope was begun below 
the fill. Unfortunately, weak market 
conditions forced suspension of opera- 
tions at the Lucky Friday Mine before the 
second stope could be completed. Al- 
though some minor rock bursting was 
encountered during development of the 
ramp system, no major design problems 
have J been found in the overall mining 
plan. 



42 



CONCLUSIONS 



The Coeur d'Alene District of northern 
Idaho is one of those mining districts 
where a unique combination of high in 
situ stresses and hard, brittle rock 
produces violent rock failures known as 
rock bursts. In the Coeur d'Alene Dis- 
trict, as in many other districts around 
the world, rock bursting did not become a 
serious operational problem until the 
mines had been developed deep into the 
rock. 

Rock bursts are violent rock failures 
that release seismic energy. When this 
radiated energy encounters a mine opening 
or is released at the face of an opening, 
the resulting rock burst may be capable 
of destroying structures or mining equip- 
ment and causing severe injuries and 
fatalities. Based on experience and the 
results of extensive studies, certain 
conditions are necessary in order for a 
seismic event to occur and release stored 
strain energy: the in situ and mining- 
induced stresses must strain some point 
or region in the rock to a condition of 
unstable equilibrium, and then an addi- 
tional stress must trigger failure. The 
failure, once initiated, must be a 
violent, brittle type of fracturing, and 
there must be available stored strain 
energy released as kinetic or seismic 
energy. The actual rock failure 



mechanisms that release seismic energy 
are only poorly understood at best, even 
after many years of intense research in 
both field and laboratory. Several 
important failure criteria have been 
postulated to explain the conditions for 
the initiation of failure, and more 
recent work has attempted to explain the 
postfailure behavior of rock. The devel- 
opment of rock behavior simulation by 
numerical modeling will provide a very 
useful means for calculating the prob- 
ability of rock bursting. 

Bureau of Mines research in rock bursts 
has focused on development of microseis- 
mic monitoring and analysis techniques, 
destress blasting and rock precondition- 
ing, and alternative stoping methods to 
reduce burst hazards. The microseismic 
method has become a refined tool for 
locating seismic events and gaining 
better understanding of rock failure 
mechanisms. In the Coeur d'Alene Mining 
District, destress blasting and rock pre- 
conditioning methods have been demon- 
strated to be effective means of working 
a burst-prone mine area. After initial 
field demonstration projects, underhand 
stoping methods are being developed fur- 
ther to permit safer mining in highly 
burst-prone, deep vein mines. 



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